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By Jeff Steinhoff

INTRODUCTION The Henderson mine utilizes a highly mechanized, continuous, panel-caving, mining system to extract ore from a deep, massive, molybdenite deposit. The mine is located 80.5 km (50 miles) west of Denver, Colorado. The mine surface facilities are located 3,170 m (10,400 feet) above sea level in a steep valley on the eastern side of the Continental Divide. Milling facilities are 24 km (15 miles) west on the western side of the divide at an elevation of 2,804 m (9,200 feet) above sea level. The ore- body is located approximately 3,000 feet south of the valley under Red Mountain. Access to the ore for men and materials is through a 915 m (3,000-foot) deep, 8.5 m (28-foot) diameter, vertical, concrete-lined, service shaft. Access from the mill is through a 15.5 km (9.6-mile) rail haulage tunnel. The mine is ventilated through an additional intake shaft and two exhaust shafts. Mine production at this time is 27,255 mtpd (30,000 stpd). The mine ventilation system supplies 1,038 cubic meters per second (2.2 million cfm) through approximately 60 miles of drifting or 2.7 tons of air per ton of ore mined. There are 130 fans in the mine in fixed locations and in vent lines with 6,900 connected horsepower in the mine. MINING METHOD AND LAYOUT The orebody is divided vertically into two major zones. The upper zone is the 8100-level production area. The bottom zone is the 7700- level production area which is in the early development stage. The rail haulage level at 7500 feet is common to both production zones. Each mine production zone consists of five associated sublevels. The cave undercut level is 16.8 m (55 feet) above the production level. Two boundary cutoff levels are located 44.2 m (145 feet) and 62.5 m (205 feet) respectively above the production level. The fresh-air level is positioned 15.2 m (50 feet) below the production level, and the exhaust vent level is 19.8 m (65 feet) below the production level. Horizontally, each production zone is divided into three panels each, 224 m (800 feet) wide. These panels are caved from south to north. As the caving in one panel nears completion, caving in the adjacent panel is initiated. Development for the caving panels is continuous so that the sublevels above the production level and the production level itself have a combination of development drifting and production-related activities. UNDERGROUND VENTILATION NETWORK The ventilation system is zoned in the same manner as the orebody itself. One major split of 600 cubic meters per second (1,270,000 cfm) ventilates the 8100-level production zone; one split of 100 cubic meters per second (210,000 cfm) ventilates the development of the 7700- level production zone; and one split of 165 cubic meters per second (350,000 cfm) ventilates the 7500 rail haulage level. The haulage tunnel requires an additional 188 cubic meters per second (400,000 cfm) of air. Development-drift ventilation is accomplished by hanging 1.0 m (3.5-foot) diameter steel ducting in the drifts with 40-horsepower, 0.96 m (38-inch) diameter fans supplying 9.4 cubic meters per second (20.000 cfm). The normal maximum length for these systems is 300 m (1,000 feet). The 8100-level production-area ventilation system is especially suited to a high level of mucking activity confined in a small area. Approximately 93 per cent of the mine's total production is transferred from the drawpoints to ore passes in 10 production drifts. The active area in each drift is 300 m (1,000 feet). Twelve 5-cubic-yard LHD units with Cat turbo- charged 170-horsepower engines are assigned to the area. Under these conditions, more than one LHD is assigned to a particular production drift. Adequate ventilation is maintained by making an air change every 97 m (320 feet) along the production drifts. Fresh air is brought into the production drifts from the fresh-air level through 1.37 m (4.5-foot) diameter raises. Air travels south along the production drift to the ore pass where it is exhausted down the ore pass to the exhaust level. The ore pass is followed by another intake which is followed by an ore-pass exhaust. At the south end of the production area, a series of exhaust fans maintain a southerly air- flow through the production level.

Jan 1, 1981

By Arthur Winslow

Location.—The Liberty Bell gold-mine is in the San Juan region, in the southwestern corner of Colorado. It is situated about 2 miles from the town of Telluride, near the summit of the Uncompahgre range, at the head of Cornett basin, at an altitude of 11,000 feet, and in a district of excellent record in the past and of great promise for the future. In the immediate vicinity are such well-known and important mines as the Smuggler Union, with a total production of over $12,000,000; the Virginius and Revenue Tunnel group, with a production of probably $5,000,000; the Tom-Boy, with a production of probably at least $3,000,000; the Japan, the Cimarron, and a number of others. History.—The first prospecting in this district is reported to

Jan 1, 1900

By R. T. Hukki

IT is no great exaggeration to say that present grinding practice and economics are largely determined by lining design. A record of outstanding liner wear can be achieved with any liner surface pattern that will positively lock the outer layer or layers of grinding media. With no slippage, lining wear is bound to be slight. At the same time, popular practice calls for tumbling loads of about 50 pct of mill volume to obtain maximum grinding capacity. Innumerable parallel grinding investigations have verified that optimum speed for such a mill lies within 70 to 85 pct of theoretical critical speed. If the mill is speeded up to 100 pct of critical, little or no grinding can be accomplished. Earlier Work on Supercritical Grinding: First investigations concerning grinding at supercritical speeds seem to be very old. Remarkable work on the subject has been performed by White,1 and his experiments have been described and summarized by Richards1 White's contributions seem to have passed unnoticed by Fahrenwald, 2,3 whose extensive experiments have been well presented, yet apparently very little appreciated. Additional work on grinding at supercritical speed has been reported, e.g., by Lewenson and Tscherny,4 USSR; Anselm5 and Grunder,6 Germany; and Rose and Evans,[6] Great Britain.

Jan 5, 1958

By L. O. Howard

MUCH work has been done recently on the lead-antimony system1 in connection with lead-rich alloys of commercial importance containing less than 20 per cent. antimony. Dean, Zickrick and Nix have called attention to the perpetuation in the literature of several erroneous data, such as a eutectic temperature of 228° C. at 12.5 per cent. antimony, instead of 247° at 13 per cent. The series is given as a classic example of a straight eutectiferous series in such late books as those of Heyn and Grossmann,2 Jeffries and Archer,3 Guillet and Portevin4 Pulsifer,5 Rosenhain,6 and Liddell,7 as well as the somewhat earlier works of Hofman (1918), Wang (1909) and Williams (1920). The work of Dean has definitely showed a solid solution of antimony in lead at about 2.5 per cent. antimony. While lecturing to a class in general metallurgy at the University of Idaho, the writer had the students prepare a series of lead-antimony alloys for the purpose of illustrating methods of thermal analysis. Proper proportions of powdered antimony (sold as C.P.) and test lead were mixed and melted in graphite crucibles under a reducing cover. Cooling

Jan 1, 1928

By L. O. Howard

Much work has been done recently on the lead-antimony system1 in connection with lead-rich alloys of commercial importance containing less than 20 per cent. antimony. Dean, Zickrick and Nix have called attention to the perpetuation in the literature of several erroneous data, such as a eutectic temperature of 228° C. at 12.5 per cent. antimony, instead of 247° at 13 per cent. The series is given as a classic example of a straight eutectiferous series in such late books as those of Heyn and Grossmann, 2 Jeffries and Archer,3 Guillet and Portevin,4 Pulsifer,5 Rosenhain,6 and Liddell,' as well as the somewhat earlier works of Hof-man (1918), Wang (1909), and Williams (1920). The work of Dean has definitely shown a solid solution of antimony in lead at about 2.5 per cent. antimony. While lecturing to a class in general metallurgy at the University of Idaho, the writer had the students prepare a series of lead-antimony alloys for the purpose of illustrating methods of thermal analysis. Proper proportions of powdered antimony (sold as C.P.) and test lead were mixed and melted in graphite crucibles under a reducing cover. Cooling

The United States Coast and Geodetic Survey has announced the completion of a new outline map of the United States on the Lambert conformal conic projection, scale 1-5,000,000. This map is intended merely as a base to which may he added any kind of special information desired. The shore line is compiled from the most recent Coast and Geodetic Survey charts, state names and boundaries, principal rivers, capitals, and the largest cities in the different states are also embodied. The value of the new outline map on the Lambert projection can best be realized when it is stated that it shows that throughout the largest and most important part of the United States, that is, between latitudes 30 ½ and 49°, the maximum scale error is only 0.5 per cent. This amount of scale error of 0.5 per cent. is frequently less than the distortion due to the method of printing and to changes from the humidity of the air. Only in southernmost Florida and Texas does this projection attain its maximum error of 233 per cent. The map measures 25 inches by 39 inches and will be sold by the Government at 25 cents.

Jan 7, 1919

By L. O. Howard

Much work has been done recently on the lead-antimony system' in connection with lead-rich alloys of commercial importance containing less than 20 per cent. antimony. Dean, Zickrick and Nix have called attention to the perpetuation in the literature of several erroneous data, such as a eutectic temperature of 228° C. at 12.5 per cent. antimony, instead of 247° at 13 per cent. The series is given as a classic example of a straight eutectiferous series in such late books as those of Heyn and Grossmann,2 Jeffries and Archer,3 Guillet and Portevin,4 Pulsifer,5 Rosenhain,6 and Liddell,7 as well as the somewhat earlier works of Hofman (1918), Wang (1909), and Williams (1920). The work of Dean has definitely shown a solid solution of antimony in lead at about 2.5 per cent. antimony. While lecturing to a class in general metallurgy at the University of Idaho, the writer had the students prepare a series of lead-antimony alloys for the purpose of illustrating methods of thermal analysis. Proper proportions of powdered antimony (sold as C.P.) and. test lead were mixed and melted in graphite crucibles under a reducing cover. Cooling

Jan 1, 1928

By W. F. Holbrook

IN a previous paper1 a method for the measurement of the comparative desulphurizing power of slags was described and data were presented covering the range of likely slags containing up to 10 per cent MgO In the present work the range is extended to include slags containing up to 20 per cent MgO, and the effect of temperature on 5 per cent and 15 pen cent MgO slags over the range 1475° to 1525° C is included The method used was that described in the previous paper except that some of the temperature measurements were made with a noble-metal thermocouple instead of a tungsten-graphite thermocouple Twelve-gram charges of 0 8 per cent sulphur-metal and 5-gram charge-, of slag were subjected to heating for 30 nun at a prescribed temperature The charges were analyzed for sulphur content For purposes of comparison, desulphurizing power is defined as the ratio of the percentage of sulphur found in the slag to that found in the metal Six slag and metal charges were made coincidentally in 5/8-in holes in a section of an electrode graphite rod One slag and metal charge was used as a control by means of which adjustment in the desulphurization value could be made for small variations in temperature and other factors For convenient presentation of the data, the slags are arranged in groups in each of which one of the four main slag components is constant Thus, the 15 per cent MgO slags are presented separately from those containing 20 per cent MgO SLAGS WITH MAGNESIUM OXIDE CONSTANT 15 Per cent MgO -The data on slags containing 15 per cent MgO cover a composition range of 24 to 50 per cent SI02, 5 to 27 per cent A1203 and 24 to 53 per cent CaO Variation in composition has been in steps of 2 per cent over much of the area Because the change in desulphurization power is generally gradual with gradual change in composition, certain sections have been covered m steps of 4I per cent and even

Jan 1, 1938

By J. E. Lawver, J. D. Raulerson, Charles C. Cook

The agriculture industry has made great strides during the past decade to increase agriculture yields through increased use of fertilizers. Increased use of fertilizers may prevent, or at least delay, mass starvation due to the alarming increase in world population. Phosphate was added to soil as a plant nutrient in the form of calcined bones at least 2000 years ago (Anon., 1964), and man has used phosphate minerals as a source of fertilization in one form or another for at least 100 years. During 1977 the world produced about 116 Mt of phosphate rock, with about 86% used for fertilizers and another 4% for animal feed supplements. More than three-fourths of the total production comes from the United States, Morocco, and the Soviet Union. From a mineral beneficiation point of view, the major sources of phosphate rock and the methods of beneficiation can be classified as follows: marine deposits not containing appreciable carbonate minerals, marine deposits requiring a francolite carbonate mineral separation, igneous deposits not containing appreciable carbonate minerals, and igneous deposits requiring apatite carbonate mineral separation. [ ] Guano, mostly from Chile and Peru, accounts for 0.1% of the total world production, and the calcium phosphates from Ocean, Nauru, and Christmas Islands and the aluminum and iron phosphates from Brazil and Aruba account for less than 4% of the world production and are thus not considered in this classification (Lawver, et al.). At present, marine phosphorite deposits account for about 75% of the world's production; the igneous deposits account for 20%. The igneous deposits low in carbonate minerals are easily concentrated by crushing, grinding, and apatite flotation. The most important igneous deposits are those of the Kola Peninsula, USSR (Woodrooffe, 1972). The igneous deposits high in carbonate materials are of corn appreciably more difficult to beneficiate, but they have been concentrated by froth flotation for a number of years. An interesting but rather complicated flowsheet of this type is at Phalabonva, in the Republic of South Africa (Lovell, 1976). The Phalaborwa deposit is an igneous complex of pyroxenite with a central core of carbonatite surrounded by a serpentine- magnetite-apatite rock called phoscorite. The phoscorite containing about 10% P2O5, 35% magnetite, and 35% calcium magnesium carbonate is currently being processed. The process involves comminuting the material for fiberation and subjecting it to a copper float using a potassium amyl xanthate as collector and triethoxybutane as a frother followed by a magnetic separation of the tailings to produce a feed for phosphate flotation. This process produces a phosphate concentrate containing greater than 36% P2O5 at a P2O5 recovery ranging from 75 to 80%. Considerable success has been claimed for recovering apatite from carbonate-bearing ores at the Jacupiranga Mine of Serrana S/A (Silva and Andery, 1972). The carbonatite currently being mined contains an average of only 5% P205 and is concentrated using a unique flotation process (Andery, 1968) to yield 96% P205 concentrates. The ore contains about 12% apatite, 5% magnetite, 80% calcite plus dolomite, and minor amounts of phlogopite, olivine, zircon, ilmenite, and pyrochlore. Feed preparation consists of crushing to -31.75 mm (-1 M in.), rod milling in closed circuit with hydrocyclones to about 92% (-50 mesh), and two-stage cyclone desliming of the -50 mesh sands at 20 m. Weight recovery in the deslimed feed is normally 85 to 88% and the corresponding P2O5 recovery is usually about 90%. The deslimed feed is conditioned at 60 to 70% solids for 15 min at pH = 8-10 with 0.6 kg/t of causticized starch for iron oxide and calcite-dolomite depression. The conditioned slurry is diluted to 20 to 30% solids, about 0.2 kg/t of fatty acid or soap collector is added to the conditioner discharge, and the reagentized ore is subjected to rougher-scavenger flotation with additional fatty acid added to the scavenger float. The scavenger concentrate is returned to rougher circuit distributor, and the rougher concentrate froth is subjected to two stages of cleaner flotation to yield a final apatite concentrate analyzing 36 to 38% P205. Flotation recovery of P205 is, in general, above 90% when treating fresh carbonatite. The high-carbonate flotation tails normally analyze 1 % P2O5 or less and are suitable for portland cement production. The marine deposits. Types 1 and 2 of central Florida are representative of enormous reserves of phosphate rock that will undoubtedly account for much of the world's production in the near future. Until very recently the sedimentary deposits high in carbonate minerals (Type 2) have not been considered reserves due to the difficulty in making a francolite-carbonate separation. Although no commercial plant has yet been built to beneficiate Type 2 ore, laboratory and pilot plant data indicate the process is viable. If so, the reserves of Florida and similar deposits throughout the world will be substantially increased. A discussion of the beneficiation of these two types of sedimentary deposits and the relation of the resulting concentrates to the fertilizer industry of the United States is the subject of this paper.

Jan 1, 1981

By A. J. Opinsky, J. L. Orehotsky, L. L. Seigle

TUNGSTEN incandescent lamp filaments possess a typical structure of elongated crystals generated upon heating the silica-alumina doped wire rapidly to 2200°C or above.&apos; It is known that these very long grains exhibit a preferred orientation, but there is disagreement in the literature about the direction. Rosi,2 and later Rieck; reported that the <135> direction of the large grains in doped wire tends to be parallel to the wire axis, but other investigators have reported that <110> is the preferred direction. Some further investigation of the orientations of large grains in tungsten wire is reported in the following. In the first series of experiments, 6-in. lengths of wire were heated rapidly in a bell jar. Current was passed through the wires and within 15 sec the wires had reached the grain coarsening temperature. 2 min was the annealing time. Wires were heated in hydrogen, nitrogen, and vacuum. Later, wires were heated in an evacuated tube furnace in which the tube was brought from 1000°C to the operating temperature within 25 sec and held 5 min at temperature. Temperatures were measured with an optical pyrometer. After heating, the wires were electro polished and etched, and the orientations of the large grains determined by the Laue back-reflection technique. Most of the experiments were carried out on a single spool of 5-mil silica-alumina doped wire, but a few tests were made with 5-mil thoriated tungsten wire and 10-mil undoped wire. 5-mil silica-alumina doped wires were reduced to smaller diameters by electropolishing for one series of experiments. Orientations of the large grains in silica-alumina doped wire heated in various atmospheres are presented in Figs. 1 and 2 as plots of the direction of the wire axis relative to the cube axes of the crystals. For most of the conditions studied, the preferred orientation cannot be described as a single direction, but the points might be said to group themselves about the arc passing through <110> and <113>. This arc is the trace of the (211) planes and contains the <135> and <124 as well as <011> directions previously observed by other investigators. The preferred orientation of large crystals in silica-alumina doped tungsten wire might be most generally described, therefore, by stating that a (211 ) tends to be parallel to the wire axis. This is true also for wires reduced in diameter by etching and furnace heated, Fig. 2, i.e., the recrystallization texture is apparently unchanged upon removal of the outer layers and substantially independent of the temperature gradient within the wire.

Jan 1, 1962

By H. R. Ogden

The notch sensitivity of titanium alloys is affected by impurity content, microstructure, and heat treatment. Using notch tensile properties to evaluate notch sensitivity, three commercial titanium-base alloys were investigated over the temperature range of -320o to 390°F. The alloys studied were Ti-5Al-2.5Sn, Ti-6Al-4V, and Ti-2Fe-2Cr-2Mo. Six interstitial levels were studied for each alloy in four mi-crostmctural conditions. The ff- 0 alloys were studied in both the annealed and heat-treated condition. All tests were conducted using 1/4-in. diam round tensile samples. The notched samples had a 50 pct notch with a notch sharpness (a/r) of 53, corresponding to Kt 6.3. Notch sensitivity was increased by increased interstitial content, low temperature, and acicular-type microstructures. The heat-treated, high-interstitial content - 13 alloys were more notch sensitive than the same alloys in the annealed condition. In addition, minima in the notch tensile-temperature curves were observed at temperatures near 32°F. These minima became more pronounced with increasing interstitial content. 1 HE increaseduse of titanium alloys in high strength applications at subzero temperatures requires that a better under standing of the low-temperature behavior of the alloys be obtained. At low temperatures, metals generally exhibit increased strength with lowered ductility and tohness. Notch sensitivity and the factors which influence it become more important as the strength level is increased and the temperature is decreased. Some of the factor s which affect notch sensitivity in ti- tanium alloys are composition, interstitial content, microstructure, and heat-treatment condition. In this paper, the results of a study of these variables on the notch sensitivity of three titanium alloys at temperatures from —320o to 390o are reported. The notch geometry selected for this program was chosen to provide a severe test of notch sensitivity. The ability of the metal to flow without initiating fracture under these severe conditions is reflected in the notch-strength values. Since notch strength is, in part, determined by the constraint exerted by the combined tensile stresses, the behavior of notch specimens reflects the properties of the material under combined stresses. For example, a material with a high notch-unnotch strength ratio may be expected to withstand the biaxial stresses encountered in a pressure vessel better than a material with a low ratio of notch-unnotch strengths. This implies that maximum strength alone is not a sufficient criterion for applications of this type. It further may be concluded that the various conditions that promote poor notch behavior also will affect adversely the performance of materials under combined stresses. In the alloys studied, these included low temperatures, high interstitial levels, and hasformed microstructures. EXPERIMENTAL PROCEDURES Three alloys, Ti-5Al-2.5Sn, Ti-6Al-4V, and Ti-2Fe-2Cr-2Mo, were purchased from commercial producers as 2-in.-dam rod stock. Oxygen and nitrogen were added to these alloys by remelting to obtain a variety of interstitial contents as shown in Table I. In order to Compensate for the much more potent strengthening effect of nitrogen compared with oxygen, an oxygen equivalent content is used to describe the total oxygen plus nitrogen contents studied. It has been suggested previously1 thatnitro-gen is approximately twice as effective in strengthening as oxygen, and carbon, to the extent it is in so-

Jan 1, 1962

By R. E. Miner, J. K. Jackson, T. L. Wilson

A study has been made of the transformation kintetics and morphology in gas-quenched Fe-5 pct Ni alloys. Massive ferrite or nlurtensite was found to form at temperatures in agreement with earlier results. Surface Preparation appeared to have the greatest effect in determining the type of transformation Product formed. Grain size change, subgrain morphology, and dislocation densities of massive ferrite were essen-tinlly invariant functions of cooling rate. The mar-tens~tic structures were typical of transformation in austenite grains of the order of the sample thickness. BY introducing the idea of a massive transformation in dilute iron-base alloys, Gilbert and Owen in 1962 served to stimulate a closer inspection of phase transformations in ferrous systems.&apos; A wealth of literature has followed concerned with the morphology,2-5 transformation mode,2&apos;3&apos;6-8 and thermodynamics of transformations.2,5,9 interesting controversies have resulted from the experimental observations; closer attention has been paid to the definition of massive transformations and still to be resolved is the Ms temperature of pure iron. The purpose of this note is to report some observations made in a detailed kinetic and metallographic examination of the transformations that occur in gas quenched Fe-5 pct Ni. The results are pertinent to the Ms extrapolations that have been made for pure iron and the observed substructure of massively transformed alloys. PROCEDURE The material used in this investigation was cast as a 5 lb vacuum induction melted, carbon deoxidized heat. The ingot was hot worked to a strip 0.10 in. thick, decarburized, and cold rolled to final thickness. Specimens were punched to sample shape and decarburized in wet hydrogen at 600°C. The chemistry of test specimens is indicated in Table I. Samples 1 by $ in., by thickness were resistance heated and rapidly cooled in an apparatus as illustrated in Fig. 1. Similar helium quenching apparatuses have been described previously;2,6&apos;10 the distinguishing feature of this apparatus is the nozzle geometry. Twin nozzles with three convergent-divergent, sonic ports were mounted perpendicular to the specimen for quenching purposes. The specimen temperature was monitored with an intrinsic 3 mil chromel-alumel thermocouple spot welded to the specimen and con- nected to an oscilloscope. A single switch stopped the current flow through the specimen, opened a gas flow solenoid valve, and initiated an oscilloscope sweep. Cooling rates were read directly from a photograph of the oscilloscope trace and the quenching rates quoted later were made at a point immediately prior to transformation on the trace. All quenching studies were conducted either under a vacuum of 10"-5 torr or a helium atmosphere. The sample geometry was chosen large enough to permit optical and electron microscopy to be performed on samples. The samples were electropolished for both types of examination in a perchloric-acetic solution. Electron microscopy was conducted on a HU 11A microscope operated at 100 kv. Dislocation densities were determined at the same general operating reflection (110) on foils thick enough to produce Kikuchi patterns. RESULTS AND DISCUSSION Transformation Temperature Observations and OD-tical Microscopy. The quenching results of this study are illustrated in Fig. -2(a). Samples which were treated in the as-decarburized condition transformed in a manner similar to that reported previously—a decrease in transformation temperature occurred as the quenching rate increased until a plateau region was reached. Within the scatter of the results obtained here, there was no break in the transformation temperature-quenching rate plot. This suggests no change in transformation mode. Further metallographic examination confirmed this observation. A martensitic transformation was obtained for similarly treated samples except that the surface was first electropolished. The Ms temperature was approximately 85°C below the plateau determined for the unpolished samples and cooling rates as slow as 2000°C per sec were adequate for producing marten-site. The transformation temperatures are in gen- / STRIP SAMPLE / COPPER I SAX/-/}&apos; / CONNECTOR-----A f, SON|C poRTS AC LEADS 3j ]j"I rofe[ \ Mf/yinK/iKj&apos;&apos;^ THERMOCOUPLE X-(^—iie"s*c? &apos; LEADS 1 INCH Fig. 1-—Schematic scction of gas quenching apparatus

Jan 1, 1970

By Cyril Wells, Robert F. Mehl

The two earlier papers in this series1,2 presented data on the rate of diffusion of carbon and of manganese in gamma iron. Apart from their scientific interest, these data are needed chiefly in order to treat data on carburization, on rate of formation of pearlite, and on rate of austenit-izing in a quantitative way.1&apos;2&apos;9&apos;10 In the present investigation rates of diffusion of nickel in ?-Fe over wide temperature and concentration ranges have been determined, and the effects of carbon, of im- purities, and of grain size on these rates have been studied. Claussen,3 using the data of Fry,4 has calculated the diffusion coefficient of nickel in y-Fe as 9.2 X 10-11 sq. cm. per sec. at 1200ºC., but had no data from which to calculate the variation of D with concentration, temperature, and percentage of carbon. Method and Materials Method.—Nickel was diffused along the axis of a cylindrical specimen across a concentration gradient, concentration-penetration curves were determined and D values calculated from Grube or Matano solutions.5 Details of the method are given

Jan 1, 1941

By Cyril Wells, Robert F. Mehl

The two earlier papers in this series1,2 presented data on the rate of diffusion of carbon and of manganese in gamma iron. Apart from their scientific interest, these data are needed chiefly in order to treat data on carburization, on rate of formation of pearlite, and on rate of austenit-izing in a quantitative way.1&apos;2&apos;9&apos;10 In the present investigation rates of diffusion of nickel in ?-Fe over wide temperature and concentration ranges have been determined, and the effects of carbon, of im- purities, and of grain size on these rates have been studied. Claussen,3 using the data of Fry,4 has calculated the diffusion coefficient of nickel in y-Fe as 9.2 X 10-11 sq. cm. per sec. at 1200ºC., but had no data from which to calculate the variation of D with concentration, temperature, and percentage of carbon. Method and Materials Method.—Nickel was diffused along the axis of a cylindrical specimen across a concentration gradient, concentration-penetration curves were determined and D values calculated from Grube or Matano solutions.5 Details of the method are given

Jan 1, 1941

By Harrison Souder

THE object of this paper is to bring to. the notice of engineers a safety detonating fuse by the use of which misfires in blasting may be eliminated and safety in blasting operations promoted. This new detonator is a French invention and is known as Cordeau detonant or detonating fuse and is sold under the name of Cordeau-Bickford. It consists of a lead tube 5 to 6 mm. in diameter, filled with trinitro-toluene. While applicable to all classes of mining it will appeal especially to those who have to do with deep-hole blasting in open-cut clines or quarries, or any operations where a large number of holes are to be shot at one time. Since the introduction of the deep-hole blasting in the Cornwall iron mines, much trouble has been experienced from misfires and many experiments were made to determine the best and safest method of blasting. Before the introduction of the detonating fuse it was found that to insure the best results, two or more high-power electric exploders, not less than No. 6, with specially insulated conductors, should be placed in each hole; the holes wired in parallel, and for a large number of holes the current to be supplied by a live wire to the bus or lead wires at about the middle of the bench. The voltage in general use is 110 to 220, but as our material is a magnetic iron ore and the holes are generally damp, we found it advisable to reduce this to about 30 volts to prevent short circuits. A No. 7 or No. 8 special insulated detonator was adopted. But with the best of care and attention to the minutest detail, we were not able to eliminate occasional misfires and the consequent ever present danger of accidents.

Jan 10, 1914

By H. H. Waller, D. L. Monroe, D. P. R. Smyth

ORIGINALLY slusher drift development was conventional, advancing the drift full 10x13-ft size at 6 ft per round. This proved dangerous and costly because the weak fractured rock of the orebody cannot be left raw for any length of time without its reacting with air to cause spalling. With this method of drift development ground conditions were worsened by many blasts and the length of time it was necessary for the drift to remain open. The problem was practically solved by driving 5x7-ft pilot drifts, then enlarging to 10x13-ft raw size by slashing with 50 to 60-ft diamond drill holes. Calculated explosive charges are used to avoid wall and roof shattering. Still more recently jacklegs and sectional steel replaced the diamond drills, resulting in the same work for less than half the cost, see Figs. 1 and 2. Crosscuts at right angles to the main haulage drifts, which cross through the center of the ore-body, divide the ore into 200-ft blocks. Eight slusher drifts on 50-ft centers draw from one block. The slusher floor is 11 ft above the rail, inclined at 1 pct grade, and the top is 18 ft below the undercut. The finished slusher drift is 102.5 or 120 ft long, depending on whether it is north or south of the crosscut. Drawpoints are staggered on each side at 17.5-ft centers. The completed drift is concreted and has an inclined ramp and a vent at the back. When crosscuts and gangways have been excavated and concreted and slusher stations installed, slusher drift development is carried out in 10 clear-cut stages: 1—Pilot drift is driven full length, 5x7 ft. 2—Station is excavated at half-way point in pilot drift. 3—Pilot drift is enlarged to slusher drift size by drilling. 4—Lower part of slash is blasted and mucked for floor concrete. 5—Floor is poured. 6—Inside upper half of slusher drift is blasted, mucked, and concreted. 7—Outside upper half of slusher drift is blasted, mucked, and concreted. 8—Drawpoints are excavated. 9—Drawpoints are concreted. 10—Concrete ramp and sheave protection recess are constructed at back end of slusher drift. A 5x7-ft pilot drift is driven full length on the center line of the proposed slusher drift using a jackleg and detachable bits on 7/8-in. hexagon steel. A 10-hp air-driven slusher and 36-in. scraper pull muck directly into cars in the main haulage crosscut, see Fig. 2. Operations are on a three-shift basis. One man mucks the previous round, drills, and blasts his round in a shift. Average advance is 5 ft per shift. When the pilot drift is completed, 7 ft of drift in the center section is enlarged to full raw slusher drift size of 10x13 ft, for the set-up to drill long slash holes. The pilot drift is slashed to full raw size from this station with jacklegs and tungsten carbide chisel bits on 7/8-in. sectional steel. Crews line the holes without trouble, working from controls established by the engineers, and the pattern of holes for the slash is arranged to give the least shattering effect

Jan 1, 1955

By Charles H. Thurman

THE data given herein were first included in a paper read before San Francisco Section, A.I.M.E., in October 1940, and are applicable to conditions existing until the gold-dredging industry was temporarily suspended by Order L-208 of the War Production Board in the fall of 1942. There is no doubt that gold dredging will be resumed at some future time, and on a scale equivalent to that of the recent past, but the cost data given herein will be affected to some extent by conditions existing at that time, and by labor rates and prices for materials then prevailing. DRAGLINE VS. BUCKET-LINE Dragline dredges should not be considered for mining large areas of low-grade placer gravel. In such deposits a dragline operator with a 5-cu. yd. dragline would be doing well to maintain an operating cost of 6¢ per cu. yd., while a bucket-line dredge with buckets no smaller than 7 cu. ft. each could be operated for less than 4¢ per cu. yd. By operating costs, I mean only field charges such as direct operating and overhead, which includes pay roll, power or fuel, repairs and maintenance, insurance, land clearing and general expenses, but not including such items as prospecting, depletion, depreciation of equipment, State Franchise Tax, or Federal Income Tax. It is not to be inferred that a dragline dredge cannot make a profit in small low-grade deposits, where the extent of a given property is too small to justify the use of the more costly bucket-line dredge. Consider small deposits of 10¢ gravel. It is obvious that a deposit of 5,000,000 cu. yd. of such gravel could not be mined profitably with a new, 7-cu. ft. bucket-line dredge, costing upward of $250,000, and with operation costs at 4¢ per cu. yd., paying for land or royalties out of earnings; nor could a new combination 5-cu. yd. dragline and washing plant, worth upward of $125,000 and working at 6¢ per cu. yd., make a profit. These comparisons are made because the hourly capacity of both types of dredges is nearly the same. The bucket-line dredge averages 93 per cent operating time in 24 hr., as against about 80 per cent or less for the dragline dredge. These percentages of operating time are important factors in making a comparison of this kind. The bucket-line dredge has greater efficiency in cleaning soft bedrock, usually characteristic of dredging deposits, and a 7-cu. ft. bucket-line dredge can be efficient to a total digging depth of 80 ft. (65 ft. below water, 15 ft. above). A 5-cu. yd. dragline dredge is limited to about 30 ft. of efficient digging depth, 5 ft. of which could be above water. In small areas of gravel not over 30 ft. deep, a 5-cu. yd. dragline dredge could make a profit where a bucket-line dredge could not do so. As an example: compare the dredging of 3,000,000 cu. yd. of 10¢ gravel 28 ft. deep with bucket line, costing 4¢ per cu. yd., and with a 5-cu. yd. dragline dredge operating at 6¢ per cu. yd. Because of small yardage, new equipment is not warranted, therefore we can assume that the dredges of both types are used equipment, fully depreciated. Dismantling, mov-

Jan 1, 1945

By Charles H. Thurman

The data given herein were first included in a paper read before San Francisco Section, A.I.M.E., in October 1940, and are applicable to conditions existing until the gold-dredging industry was temporarily suspended by Order L-208 of the War Production Board in the fall of 1942. There is no doubt that gold dredging will be resumed at some future time, and on a scale equivalent to that of the recent past, but the cost data given herein will be affected to some extent by conditions existing at that time, and by labor rates and prices for materials then prevailing. Dragline vs. Bucket-line Dragline dredges should not be considered for mining large areas of low-grade placer gravel. In such deposits a dragline operator with a 5-cu. yd. dragline would be doing well to maintain an operating cost of 6¢ per cu. yd., while a bucket-line dredge with buckets no smaller than 7 cu. ft. each could be operated for less than 4¢ per cu. yd. By operating costs, I mean only field charges such as direct operating and overhead, which includes pay roll, power or fuel, repairs and maintenance, insurance, land clearing and general expenses, but not including such items as prospecting, depletion, depreciation of equipment, State Franchise Tax, or Federal Income Tax. It is not to be inferred that a dragline dredge cannot make a profit in small low-grade deposits, where the extent of a given property is too small to justify the use of the more costly bucket-line dredge. Consider small deposits of 106 gravel. It is obvious that a deposit of 5,000,000 cu. yd. of such gravel could not be mined profitably with a new 7-cu. ft, bucket-line dredge, costing upward of $250,000, and in operation costs at 4¢ per cu. yd., paying for land or royalties out of earnings; nor could a new combination 5-cu. yd. dragline and washing plant, worth upward of $125,000 and working at 6¢ Per cu. yd., make a profit. These comparisons are made because the hourly capacity of both types of dredges is nearly the same. The bucket-line dredge averages 93 per cent operating time in 24 hr., as against about 80 per cent or less for the dragline dredge. These Percentages of operating time are important factors in making a comparison of this kind- The bucket-line dredge has greater efficiency in cleaning soft bedrock, usually characteristic of dredging deposits, and a 7-cu. ft. bucket-line dredge can be efficient to a total digging depth of 80 ft- (65 ft. below water, 15 ft. above). A 5-cu. Yd. dragline dredge is limited to about 30 ft. of efficient digging depth, 5 ft. of which could be above water. In small areas of gravel not over 30 ft. deep, a 5-cu. yd. dragline dredge could make a profit where a bucket-line dredge could not do so. As an example: compare the dredging of 3,000,000 cu. yd. of 10 gravel 28 ft. deep with bucket line, costin 4¢ Per cu. yd., and with a 5-cu. yd. dragline dredge operating at 6¢ Per cu. yd. Because of small yardage, new equipment is not warranted, therefore we can assume that the dredges of both types are used equip-ment, fully depreciated. Dismantling, mov-

Jan 1, 1946

H. W. Turner, Portland, Ore. (communication to the Secretary*) :—Mr. Brown&apos;s paper on the complicated groups of veins exposed in the workings of the Bodie mine forme a valuable contribution to the literature of the structural features of ore-depoeits. It is much to be desired that managers should put into print the results of their observations of the mines under their charge. These observations, as well as data of production, kind of ores, etc., usually accumulate for years, and unless the mining geologists of the Geological Survey happen along before the mine is worked out, such data seldom get into print. About the year 1899 I went through the mines at Bodie with R. C. Turner, then superintendent, and made some notes on the veins and gangue-minerals, and collected some specimens. Mr. Brown refers to crystals of feldspar (albite) found on fragments of quartz on the surface. I collected a number of specimens of these crystals, mostly from Queen Bee hill, and

Jan 1, 1909

By Arthur F. Taggart

Tins paper proposes two units (believed to be new) for designating, respectively, capacity and efficiency for primary and intermediate crushers. CAPACITY Operators know that the tonnage of rock that can be put through a particular crusher in a give time depends primarily upon the crusher setting (open setting for jaw and primary gyratory types; closed setting for cones and rolls), and secondarily on the lithological character of the feed, its maximum size, the amount of fines that it carries, its moisture content, and the method of feeding. The proposed capacity formula states a relationship between all of these factors in one empirical equation, as follows: TR = TR80KK&apos;K" [I] in which Ts, called "reduction tons per hour" is the new capacity unit; T = tons of feed per hour requiring crushing; R80 = "80-per cent reduction ratio"; and K, K&apos;, and K" are constants that approximate quantitatively lithological character, moisture content and method of feeding, respectively. T is the hourly weight of material in the feed that is coarser than the coarsest discharge of the crusher. (For marked departures from the normal specific gravity range - 2.5 to 2.7 - T should. be adjusted to corresponding volumes.) In the crushers under consideration, discharge from the crushing zone is limited by the minimum dimension of a particle. Passage of the particle through a square-mesh screen is limited by its intermediate dimension. Sheppard1 has shown that the ratio of maximum: intermediate: minimum average dimensions of broken rock fragments may be taken as about 1.5 : 1 : 0.6 for nonslabby rocks such as are normally crushed, while for slabby rocks, such as thin-bedded sedimentaries, foliated schists and gneisses, and the blocky metamorphics, the ratio is of the average order I.3:I:0.3, with the possibility of an even smaller relative value for the thickness. Gaudin2 finds the corresponding ratio for 150 + 200-mesh pyrite to be 1.4: 1 : 0.6. The hourly tonnage of crusher-plant feed, therefore, should be diminished by the amount thereof that would pass a square-mesh screen of which the aperture is S times the crusher setting, where S is a shape factor with the value 1.7 for rocks that break exceptionally cubic; 2 for ores and rocks that exhibit no appreciable tendency to slabbiness; and 2.5 to 3, or, in exceptional cases, even larger, when the tendency to slabbiness is marked. R80 is the quotient of the theoretical square-mesh aperture that would pass 80 per cent of the feed, divided by the aperture that would pass 80 per cent of the product. It is best determined from screen tests.

Jan 1, 1941